Process of recovering precious metals from refractory source materials



United States '7 atent 2,777,764 Fatented Jan. 15, 1957 Norman Hadleyand Howard Tabachnick, Stamford,

Comm, assignors to American Cyanamid (Iompany, New York, N. v acorporation of Maine No Drawing. Application July 9, 1954,

Serial No. 442,454 19 Claims. (Cl. 75-105 This invention relates to aprocess for the recovery of locked precious metals from refractorysources, More specifically, it relates to a process of recovering lockedprecious metals from refractory sources which comprises slurrying saidrefractory source of said precious metals in a non-alkaline aqueousmedium, heat' g said slurry under pressure to l20-330 C. in the presenceof an oxidizing agent, separating the undissolved matter and recoveringthe precious metal in the undissolved matter by cyanidation.

The recovery of precious metals from ores by cyanide.- tion is a processknown in the art. It is used widely, especially in the recovery of goldand silver. While it is -a very effective method of recovering theseprecious metals, it has been known for a long time that with cer-. taintypes of ores, this method is not usable. In these types of ores, theprecious metal is occluded in another; mineral in such a manner that thecyanide solution cannot contact the precious metal to dissolve it. Suchoceluded precious metals are said to be locked and these ores are spokenof as refractory ores. These refractoryores are those which containsulfur, arsenic, selenium, antimony, tellurium or carbonaceous matter.They include such important minerals as pyrite and arsenopyrite.

nides, antimony compounds, tellurides and the like, into oxides. Some ofthese oxides are volatile and escape the reaction mixture 1n thismanner. Others are now porous and more easily penetrable, permitting thecyanide to attack the precious metal.

Evenv this calcining process has strong disadvantages. First of all,thetemperature must be controlled rather carefully. If the temperaturegets too high, the. oxides which are being formed tend to fuse andreoccludethe precious metal. If the temperatureis not high enough, thereis incomplete conversion of the locked ore into an oxide orewithresulting loss of precious metalstill-locked therein.

This method has been used over a long period, asthe. only way in whichthis type of ore could'be attacked. However, the difficulties enumeratedabove have resulted in tailings from such workings which still containsignifi; cant amounts of precious metals. accumulated in largequantities in a great many mining operations and have been held in thehope that a good method of working them up would be discovered. Such x dzin re t, pper, c, bait. nd i l an the like, under pressure and hightemperature, in ordejr thee to dissolve these base metals without needof smelting. However, such a process will not result in getting anyprecious metals present in solution. it is one of the advantages of ourinvention that we have found how this new metallurgical technique can beused to great advantage on sources of precious metals which methods ofthe prior art could not satisfactorily work up.

We have found that it is possible to recover precious sists inpre-treating the said locked source material for precious metal in anon-alkaline oxidizing medium under pressure, at a high temperature,separating the undis solved matter, and recovering the precious metal inthe undissolved matter by various customary means, such as cyanidation.

This method is operable with any precious metal ore, but wherecyanidation will directly attack the precious metal it is often noteconomic. Refractory sources in which the precious metal is locked oroccluded, however, can be worked up satisfactorily only by thistreatment, and the present invention is particularly advantageous withthese sources. Such refractory sources consist of, for example, ores orconcentrates in which the precious metals are intimately associated withpyrite and arseno- 1 by any of the usual methods.

alkaline.

a pose ore sulfides.

The aqueous medium in which the pressure oxid'aticin of' the refractorysource material for precious minerals is carried" out is described inthis specification asnonthe beginning of the reaction. The acid can beany moderate or strong mineral acid which tends to decom Because of itsch'eapness and availability, sulfuric acid is preferred. However,nitric, phos phoric, hydrochloric, hydrobromic and similar acids areequally usable. Strong organic acids such as trichloracetic acid and thelike, especially when they are somewhat tion, but because of the highercost, they are not so economically useful. It is preferred thatbetween 1. and 5% acid: solution be used as the aqueous medium for thepressure oxidation.

oxidizing agents as hydrogen peroxide, alkali metal per manganate,sodium dichromate and the like, the cost of these oxidizing agents makesthem less attractive. By far the cheapest andpreferred oxidizing agentisair or compressed oxygen and these form the preferred species to beused. It is desirable. that the pressure reaction be carried out in astrongly oxidizing medium and that sufficient oxidizing agent besupplied to oxidize relatively completely all the minerals which areoccluding "the precious metals. The nature of the oxidizing agent is farless important than that it be present in sufficient' quantity. Theautoclaving is carried out above 120 C. Higher temperatures providespeedier oxidation and therefore consume much less oxidizing agent. Itis therefore'preferred, usually, to carry out the autoclaying at a muchmore elevated temperature, such as about 330 C. The'pressure may be thevapor pressure of water at the temperature used or further pressure canbe supplied by adding oxygen orair under pressure. The reaction occursbestif pressures over 100 p. s. i. g. are used.

The reactions in the autoclave produce much 'acid.- I

Since cyanidation ooerationsmust be kept alkaline, the

acid produced in the autoclaving introduces a problem of limeconsumption in the subsequent cvanidation. The product of theautoclaving can sometimes be washed reasonably free of acid beforecyanidation, and this can keep the 'alkalizing problem to a minimum. Insuch cases high tem eratures such as 320 C. are referred. However, inother cases, such washing is not eflicient or else consumes so muchwater as to be quite uneconomical. In'addition, some of the insolubleproducts of the acid oxidation consume more lime than others. Apparentlythe temperature at which the autoclave reaction was carried out affectsthe alk li consuming nature of insoluble products formed. In these casesanother embodiment of our invention assumes im ortance.

, We have further found that the consumption of lime in the cvanidationrocess can be greatly reduced bv controlling the autoclaving tem eratureto betweenxl SO-and 250 C. while keeping the pressure in the preferredrange, near650 p. s. i. g., by suoolving additional oxvgenor air to theautoclave. A temperature of about 175200 C. is

preferred for arsenical concentrates with slightly hi her tem eraturesbetter for antimonial concentrates. Under such conditions it is possibleto save as much as 300 pounds of lime. per ton of concentrate. Where thenature of the locked source material is such that the autoclaved pulpcannot be efliciently or economically washed, the

use of this lowertemperature range shows great economic advantages. Thecost oflarge quantities of lime, especially when the consumer is locatedgreat distances from the source of such material. is a major factor.

The precious metals which occur most frequently in a locked form in arefractory source material are gold and silver. However. the otherprecious metals such as platinum, osmium, iridium, palladium and thelike also somei times occur in similar form and ourinvention isequallyusable with them. Thus any precious metal which'will not dissolve in an'acid-ox-idizing medium but which will dis solve on cyanidation canreadily be used in the process of our invention. The wide applicabilityof this process for the extraction of precious metals from refactoryores and similar sources can be seen from the following wide variety ofmaterials to which it can be successfully applied. The following tablegives the assay and analysis of six representative samples of this typeof material:

- cyanide solution.

residue and filtrate then being assayed for gold content,

Sample Sample 5 6 Gold, oz.lton. 0. 404

Silver, oz./ton

Iron, percent-.." 2. 94 23 49 Sulfur, percent--- 5.19 Arsenic,percent... 7. l7

Antimony, percent Nickel, percent Copper, percent Zinc, percentTellurium, per- Present -i;

cen Insolubles, percen Carbon. percent. Manganese, percent These samplescan be described as follows: Sample 1.A gold ore from Nevada containingthe minerals realgar (AS252) orpiment (AS253), pyrite (Fess) quartz, andcarbonaceous material. Sample 2.-A. flotation concentrate from Canada hgh in sulfides. The chief minerals are arsenopyrite, pynte'.

with smaller amounts of stib'nite, sphalerite, and hi kel ,7

mineral. v

' Sample 3 .-The residue left after roasting and cyaniding Sample 2.Consists mostly of ferric oxide (FezOs).

Sample 4.-A flotation concentrate from Colorado in which the goldoccurs-chiefly in the form of tellurides.

Other minerals are pyrite and quartz.

Sample 5.--An antimony flotation concentrate from Southern Rhodesia. Thechief mineral is berthierite (FcS-SbzSz), with some arsenopyrite andpyrite.

' Sample 6.--An arsenic flotation concentrate from Southern Rhodesia.with a minor amount of berthierite.

' Our invention can be described by the following examples in which theuse of our invention is compared to processes previously known and inwhich parts are by weight unless otherwise specified.

EXAMPLE 1 I A charge of Sample I was ground until 80% passed through a200 mesh screen and cyanided 24 hours at a pulp density of 25% solidsinan alkaline cyanide 501! tion. The pulp was then filtered and washed.The residue and'the solution were assayed for gold content, andthc goldextraction was calculated with the following results.

Gold extraction 66.3%

Gold in residue 0.138 oz./ton.'

Sodium cyanide consumption 15.1 lb./ ton of ore.

" iEXAMPLE 2 A charge of Sample 1 was subjected to acid oxidation in thepresence of dilute sulfuric acid, in the proportion of 1 part ore to 3parts 5% H2804. The slurry wasretained in the autoclave for 5 /2 hoursunder 840 to 1000 lb. pressure per square inch (water vapor plussupplied oxygen), and at a temperature of 230 C. to 245 C. The

slurry was then cooled, filtered, and the residuewas washed with water.The washed residuewas cyanided for 24 hours at a concentration of 25%solidsin alkaline The pulp was filtered and washed, the

with the following results: Gold extraction 91.9%. x 7 0.031 oz./ton.

' Gold in residue Sodium cyanide consumption EXAMPLE 3 A charge ofSample 2, as received, was cyanided for 72 hours in alkaline cyanidesolution at a pulp density, of 14% solids. The slurry wasfiltered,washed, and the products were assayed for gold content. From theseassays goldcxtraction wascalculated, withthe following results:

The chief mineral is arsenopyrite,

1 2,777,764 s 6 Gold extraction u ,a i 59.2%. EXAMPLE 8 g g {681d}? .fgfi pf A charge of Sample 4 was subjected to acid-oxidation o um cyam econsumptmn on in an autoclave in the presence of dilute sulfuric acid.concentrate The slurry consisted of 1- part Sample 4 to 3 parts EXAMPLE4 5 sulfuric acid The conditions were 2 hours in the autolave at 225 C.and 600 lbs. per square inch pressure A char e of Sam le 2, as received,was roasted under 6 oxidizing ionditions in a muflie furnace for 3 /2hours the (Water vapor plus supphed oxygen) The slurry-Was thenfinishing temperature being C calcifie"was cooled, filtered and washedThe washed residue was washed with water to remove soluble salts. Thewashed 10 f g il kgti gj aq tg' 202560113? ft); 4211:2112; calcine wasthen cyanided for 48 hours according to usual 9 n m cyam u n e p p aspractice The Slurry was filtered and Washed and the and washed The cyame residue and the filtrate welre residue and the solution were assayedfor gold content. i for goldl a the results the go d The amount of goldextracted by the cyanide solution ex rac Ion was ca cu a e as 0calculated, with the following results: Gold extraction 99.5%.

Gold in residue 0.03 o /ton. 33 3 $2233: g 'g ggy i m y ni n ump i n-fln 0-5 Ib- L Sodium cyanide consumption 0.81 lb./ton EXAMPLE 9concentrate go A charge of Sample 5, as received, was cyanided for X LE5 72 hours in alkaline cyanide solution at 16% solids. The lurry wasfiltered and washed. The cyanide residue and A char e of Sam le 2 wassub ected to acid-oxidation with diluti sulfuric acid. The sliirryconsisted of 1 part the 801M911 were assaye.d for gold content and fromthe Sample 2 to 3 parts 5% sulfuric acid. The tempefamm 0 assays thegold extraction was calculated, as follows. in the autoclave was 225 C.and the pressure 650 lbs. Gold extraction 55.34%. per square inch (watervapor plus supplied oxygen). Gold in residue 2.48 0z./ton. The time ofthe claving was 2 hours. The slurry was Sodium cyanide consumption 24.1lbs/ton. then filtered and washed, and the acid-oxidation residueEXAMPLE 10 was sub ected to cyanidation for 48 hours at a pulp den- 0sity of 10% sollds in an alkaline cyanide solution The P 5, as received,was wasted under pulp was filtered and washed The cyamdation residueOXldIZlfig condltlons In a muffle fumacfi for 31/2 hours, and thecyanide solution were assayed for gold content, h filllshlflg mp bemgThe Q1 W from which results gold extraction was calculated, as groundWei 111ml 60% Passed through ZQQ m fl f ll It was then filtered andwashed to remove any soluble I salts. The washed calcine was thencyanided for 48 GO d .extragnun hours at a concentration of 16% solidsaccording to the Golgi m T ""T 2 5 usual practice. The pulp was filteredand washed. The sodmm cyamde consumpnon sjton cyanide residue and thesolution were assayed for gold concentratecontent, and the goldextraction was calculated, as fol- EXAMPLE 6 4O lows;

Four parts of Sample 2 and 1 part of Sample 3 were Gold extraction35.39%. blended. Sample 3, as mentioned previously, is a residue Gold inresidue 3.30 oZ./tou. resulting from roasting and cyaniding Sample 2, i.e., in Sodium cyanide consumed 4.4 lbs/ton. its present form it containsno cyanide-soluble gold. The EXAMPLE 11 blend was slurried with dilutesulfuric acid in the pro- 6 portion 1 part blend to 3 parts 5% sulfuricacid. The A Charge of a p 5 Was sublectefi t0 f Pn slurry wastransferred to an autoclave and subjected to 111 an auto'clave m P 0fdlhlt? 1m 9 1- 650 lbs. per square inch pressure (water vapor plus sup-The Slurry f f OI P of 121 5 to 3 P t5 plied oxygen). The temperaturewas maintained at 223 of 5% sulfuric acid. ne temperature in theautoclave C. to 248 C. for 3 hours. After cooling, the slurry was was Cand the P F 659 P fl n filtered and washed. The residue was cyanided for48 (Watfif Vapor P1115 pp f Y The tlme 31 hours at 16% sch-d5 in analkaline cyanide solution The autoclave was 2 hours. the slurry was thenremoved Pulp was then filtered and Washed The residue and tha from theautoclave, filtered and washed. The ac1d:ox1dafiltrate were assayed forgold content and the percentages {1.0K Was subjicted to cyamdanonalkalme of gold extracted from Samples 2 and 3 were calculated amae501mm f 43 Pulp Was filtered and with the 011 owing results, washed, theresidue and solution being assayed for gold T 1 b 97 4? content, withthe following results:

em go d extraction from lend a. v u

. Gold in blend residue 0.095 oz./ton. 5 Fg g g-3 A Gold extraction fromSample 2 99.1%. 0 m resl F f. G01 d extracfign from Sample 3 Sodiumcyanide consumption 1.28 lb./tou. Sodium cyanide consumption 9.97lbs/ton EXAMPLE 12 of blend A charge of Sample 6, as received, wascyanided for EXAMPLE 7 72 hours in alkaline cyanide solution. The pulpwas g then filtered and washed The residue and the filtrate g g i s f fi f with an alkaline were assayed for gold content and the goldextraction I yamde solution for 1.. hours at a pulp density wascalculated as fOHOWS of 25% solids. The pulp was filtered and Washed.The cyanide residue and the filtrate were assayed for gold PO GOIQextrac tlon content and from these assays the gold extracted by Golf} mresldfle oL/ton' "cyanide solution was calculated, as follows: Sodlumcyanide m Ill/ton Gold extraction 85.8%. EXAMPLE 13 :Gold in residue0.90 oz./ton. A charge of Sample 6, as received, was roasted underSodium cyanide consumed 1.51lbs./ton. oxidizing conditions in a mufilefurnace for 3V2 hours,

. acid-oxidation Gold extraction 71.18%. A Gold in residue 0.435 Oz./tn.Sodium cyanide consumption 0.93 1b./ton.

EXAMPLE 14 A charge of Sample 6, as received, was subjected to in anautoclave in the presence of dilute sulfuric acid. The ratio of Sample 6to acid was 1:3; The autoclave conditions were 225 C. and 650 lbs. persquareinch (water vapor plus supplied oxygen). The time of contact underwas 2 hours. The slurry was cooled, filtered and washed. The residue wasthen subjected to cyanidation for 48 hours in alkaline cyanide solution.The pulp was filtered and washed. The residue and the solution wereassayed for gold content and the gold extraction calculated, as follows:

Gold extraction Gold in residue Sodium cyanide consumption 97.85%. 0.04oz./ton. 1.06 lb./ton.

EXAMPLE Sample 2 was treated by various methods to recover the gold bycyanidation. I Cyanidation of this concentrate without pretreatmentrecovered 59.20% of the gold with the consumption of 4.77 pounds per tonof NaCN equivalent of Aero brand cyanide and 13.54 pounds of CaO. I,Sample 2 was subjected to a roast and the resulting calcine cyanided.By this method of treatment 86.05%' of the gold was recovered bycyanidation. Reagent eonsumptions were 0.52 pound of NaCN equivalent ofAero brand cyanide and 5.57 pounds of CaO per ton of con-- centrate. Thecyanide residue assayed 0.82 ounce gold per ton (equivalent to 0.62ounce per ton of concentrate).

Sample 2 was pulped at 25% solids with 5% sulfuric acid solution in anautoclave and heated to 250 C. with 650 p. s. i. pressure (oxygen plusvapor). Oxygen was added to maintain the 650 p. s. i-. operatingpressure; contact time was 2 hours. The autoclave product was filteredand washed, repulped, again filtered and washed,

and then cyanided. Gold recovery. by cyanidation was' 1.54 pounds ofNaCN 99.13%; reagent consumptions were equivalent of Aero brand cyanideand 362.8 pounds of CaO per ton of concentrate.

Four additional tests clave conditions except temperature were heldconstant. These tests were completed with autoclave temperatures of 225,200, 175 and 150 C. The results of these tests are given below:

Consumption 1 NaCN CaO Recovery,

Temperapercent Pressure, ture C y 1 Piounds per ton oi concentrate.

EXAMPLE 16 Sample 6 was treated 'as Sample 2. Results of test work aregiven below:

sulfuric these conditions were completed in which all autoin the sameseries of treatments CaO consumption I approximately 12 times 8 Directcyanidation s Gold recovery 17.69%. 3

NaCN consumption -3 5.81 lb ./ton.

24.18 lb./ton.

Roast-cyanidation of calcine 71.18 0.93 lb./ ton. 14.77 lb./ton.

Gold recovery u NaCN consumption CaO consumption Acidoxidatiofl-cyanidation At 250 centigrade:

Gold recovery 97.85%. NaCN consumption 1.06 1b./ton. CaO consumption..2:66.2lb./t'on.

At 175 centigrade:

Gold recovery 95.14%. j NaCN consumption 0.83 lb./ton. CaO consumption30.0 lb./ton.

EXAMPLE 17 Sample 5 was treated in a manner similar to that of Example15. Results of test work on this product are given below:

' Direct cyanidation V 55.34%. 24.12 lb./ton. 23.7 lb./ton.

Gold recovery NaCN consumption CaO consumption Roast-'cyanidation ofcalcine Gold recovery 35.39%. NaCNconsumption 4.42 lb./ton. CaOconsumption 2.23 -lb./ ton.

Acid oxidation-cyanidation Consumption 1 Tenipera- Pressure, Recovery,

ture 0. p. s. i. percent NaON CaO V V 175 650 27. 69 48. 5 21. 200 G 21.05 30. 4 50. 94 225 650 7. 62 26. 9 90. 94 250 65 0 1. 08 98. 3 97. 57

1 Pounds per ton of concentrate.

The mine from which Samples 5 and 6'came produces as much arsenicalconcentrate as antimony concentrate. When the two concentrates weremixed in that proportion and autoclaved at C. relative completedestruction or the sulfide minerals was obtained with much lower limeconsumption at the lower temperature. Results of these two tests aregiven below;

' Consumption 1 Test No. Tempera- Pressure, Recovery,

ture C. vp. s; 1. percent NaCN 09.0 V

1 Pounds per ton 0t concentrate.

4. The process according to claim 3 in which the gaseous oxygen ispresent as air.

5. A process of recovering precious metals chosen from the groupconsisting silver and gold from refractory source materials, whichcomprises heating a slurry of said refractory source material in anaqueous acid medium in the presence of gaseous oxygen to a temperatureover 120 C., recovering from the said slurry the insoluble material, andrecovering the precious metals from the said insoluble material bycyanidation.

6. The process according to claim 5 in which the refractory sourcematerial is an ore containing arsenic.

7. The process according to claim 6 in which the source materialcontains arsenopyrite.

8. The process according to claim 5 in which the source material is asulfide ore.

9. The process according to claim 8 in which the source materialcontains pyrite.

10. The process according to claim 5 in which the source material is acalcined ore.

11. The process according to claim 5 in which the source materialcontains tellurides.

12. The process according to claim 5 source material contains antimony.

13. A process of recovering locked gold from arsenic containing ores,which comprises heating said ore in an in which the aqueous acid mediumin the presence of to a temperature of approximately 330 from the saidslurry the gaseous oxyge: C., recoverin; insoluble material, and recyanidation.

14. The process according to claim 8 precious metal is gold andapproximately 330 C.

in which the the temperature of heating is References Cited in the fileof this patent UNITED STATES PATENTS 1,648,760 Dietzsch Nov. 8, 19272,686,114 McGauley et a1. Aug. 10, 1954 FOREIGN PATENTS 19,171 GreatBritain of 1899

1. A PROCESS OF RECOVERING LOCK PRECIOUS METALS FROM REFRACTORY SOURCEMATERIALS WHICH COMPRISES HEATING UNDER PRESSURE SAID SOURCE MATERIAL INA SLURRY IN AN AQUEOUS NON-ALKALINE OXIDIZING MDEIUM TO A TEMPERATUREOVER 120*C., RECOVERING FROM THE SAID SLURRY THE INSOLUBLE MATERIAL, ANDRECOVERING THE PRECIOUS METALS FROM SAID INSOLUBLE MATERIAL BYCYANIDATION.